Separation of titanium halides from aqueous solutions

ABSTRACT

A method for the production of titanium metal from titanium-bearing ore. The method comprises leaching said ore or a concentrate thereof with an aqueous solution of a hydrogen halide; separating solids from the leach solution, to provide a leachate solution. The leachate solution may be subjected to extraction with an immiscible organic phase. Titanium halide is separated from the organic phase by stripping. Preferably, the titanium halide is titanium tetrachloride.

FIELD OF THE INVENTION

The present invention relates to methods for the separation of titaniumfrom ore, especially iron-bearing ore e.g. ilmenite ore. In embodimentsof the invention, the method relates to the recovery of titaniumtetrahalides, especially titanium tetrachloride from solutions. Infurther embodiments, the invention relates to recovery of titanium metalfrom such ore.

BACKGROUND OF THE INVENTION

Many processes are known for the recovery of titanium dioxide from ores.Ilmenite, which contains mainly titanium oxide and iron oxide values,often is employed in such processes. The majority of processes for therecovery of titanium dioxide from ores involve digestion of the ore in amineral acid, such as hydrochloric acid or sulphuric acid, to remove atleast the titanium values from the ore. In such processes, however, thepurity of the titanium dioxide produced is about 90-95%, and hencefurther purification procedures may be required to produce a pigmentgrade product, which adds considerably to the cost. Many of the furtherpurification procedures involve techniques that are environmentallyunacceptable without extensive procedures to treat various solutions andsolids obtained. Such treatment processes tend to be costly.

Processes for the recovery of titanium dioxide from ilmenite in highpurity and high yield are known. One such process is described in U.S.Pat. No. 3,903,239 of S. A. Berkovich, which discloses a process whichcomprises contacting ilmenite, or a concentrate thereof, in particulateform with concentrated hydrochloric acid at a temperature of about15-30° C. to solubilize and leach from the ore at least 80%, preferablyat least 95%, of the iron and titanium values. The leaching operationmay be carried out over an extended period of time, typically from 3-25days, depending on the technique employed and the quantity of ironvalues to be recovered. Leaching techniques include counter-current flowor the use of closed cycle loops in which hydrochloric acid iscontinuously passed through a bed of the ore. The leaching operation isexothermic, and the reactants are maintained in a temperature range of15-30° C. by cooling, if necessary.

The ilmenite or similar ore used in the process may be treated as suchor may be beneficiated to form a concentrate in any desired manner.Ilmenite generally contains TiO₂.FeO with varying amounts of Fe₂O₃ andgangue materials, usually silicates, alumina, lime and magnesium.Beneficiation may be employed when the ore is of low TiO₂ content.

The ore or concentrate may be pre-treated prior to contact with theconcentrated hydrochloric acid to increase the rate of dissolution ofthe titanium and iron values during the leaching step. Suchpre-treatment may include an initial oxidation at elevated temperature,such as from 600-1000° C., in the presence of air and/or oxygen to splitthe TiO₂.FeO followed by reduction of at least part of the iron oxidewith carbon or carbon monoxide. This is a smelting step, with slag fromthe smelting step being fed to the leaching step and pig iron beingmarketed.

Subsequent to the leaching step, it is necessary to convert any ferriciron in the solution to ferrous iron, which is typically achieved byreduction of the ferric iron in the leach liquor with a gaseous reducingagent e.g. sulphur dioxide. The conversion of ferric iron to ferrousiron in this manner is essential in view of the affinity of titaniumdioxide for ferric iron and the difficulty in separating ferric ironfrom titanium dioxide.

The solution of titanium chlorides and ferrous chloride which is thusobtained, and which may contain minor quantities of gangue metalchlorides, typically calcium and magnesium materials, is then mixed withwater to cause hydrolysis of the titanium chlorides. A seeding amount,generally about 1-2%, by weight of the titanium oxyhydrate to beprecipitated (TiO_(2.3)H₂O) is included in the mixture. Titaniumoxyhydrate precipitates from the mixture. The hydrolysis is carried outusing a quantity of water at least sufficient to precipitatesubstantially all of the titanium values from the solution butinsufficient to cause precipitation of other metal oxides or hydroxides.The titanium oxyhydrate that is precipitated from the mother liquor isthen washed substantially free of entrained mother liquor and dried. Thewashed precipitate is converted at elevated temperature, typically700-1000° C., in the presence of air and/or oxygen into the anatase orrutile form of titanium dioxide.

Alternative methods that require less treatment of solution and solidsto ensure environmental acceptance and/or are less expensive inachieving environmental acceptance, as well as producing titanium andother products of high value e.g. high purity, are required.

SUMMARY OF THE INVENTION

A method for the separation of titanium, and for production of titaniummetal, from titanium-bearing ore that involves a reduced number of stepshas now been found.

Accordingly, one aspect of the present invention provides a method forthe separation of iron values from titanium-bearing ore, comprising thesteps of:

a) leaching said ore, or a concentrate thereof, with an aqueous solutionof a hydrogen halide;

b) separating solids from the leach solution obtained in (a), to providea leachate solution;

c) subjecting the leachate solution to extraction with an immiscibleorganic phase that selectively extracts iron values into said organicphase, titanium values in the leachate solution selectively remaining inthe aqueous leachate solution.

Another aspect of the invention provides a method for the production oftitanium metal from titanium-bearing ore, comprising the steps of:

a) leaching said ore or a concentrate thereof with an aqueous solutionof a hydrogen halide;

b) separating solids from the leach solution obtained in (a), to providea leachate solution;

c) subjecting the leachate solution to extraction with an immiscibleorganic phase having a boiling point that differs from the boiling pointof the titanium halide in the leachate by an amount that permitsseparation thereof by fractional distillation, said organic phase beingstable with respect to the titanium halide; and

d) stripping titanium halide from the organic phase obtained in step (c)by heating to volatilize the titanium halide and effect separation fromthe organic phase.

A further aspect of the invention provides a method of separating atitanium halide from a concentrated aqueous solution of the titaniumhalide, said titanium halide being in a concentration such that thetitanium halide is substantially stable in said aqueous solution,comprising:

a) admixing said aqueous solution with an organic phase having a boilingpoint that differs from the boiling point of the titanium halide by anamount that permits separation thereof by fractional distillation;

b) separating the organic phase so obtained from the aqueous solution;and

c) heating the organic phase and stripping titanium halide therefrom.

Yet another aspect of the invention provides a method for the separationof titanium from a titanium-bearing ore, said ore containing iron,comprising the steps of:

a) leaching said ore, or a concentrate thereof, with an aqueous solutionof a hydrogen halide in the presence of an oxidising agent; and

b) effecting a separation of titanium dioxide obtained in step (a) fromsaid solution.

A further aspect of the invention provides a method of forming atitania-rich slag from a titanium-bearing ore that contains iron,comprising the steps of:

a) calcining the ore under oxidizing conditions to eliminate sulphurfrom said ore, said calcining being carried out at a temperature of atleast 1200° C.;

b) subjecting the hot calcined ore of step (a) to reducing conditions inthe presence of CO;

c) transferring the hot reduced calcined ore obtained in step (b) to asmelting step;

controlling the reducing conditions and the smelting step to obtain pigiron and a titania-rich slag with a predetermined iron content.

Another aspect of the invention provides a method of forming atitania-rich slag from a titanium-bearing ore that contains iron,comprising the steps of:

a) calcining the ore under oxidizing conditions to eliminate sulphurfrom said ore, said calcining being carried out at a temperature of atleast 1200° C.;

b) subjecting the hot calcined ore of step (a) to reducing conditions inthe presence of CO;

c) transferring the reduced calcined ore obtained in step (b) to aleaching step in an aqueous solution of weak sulphuric acid; controllingthe reducing and leaching conditions to obtain a titania-rich materialhaving less than 5% by weight of the iron content of the ore.

A further aspect of the invention provides a method for the separationof iron and titanium values from an iron/titanium ore, comprising thesteps of:

a) leaching said ore, or a concentrate thereof, with an aqueous solutionof a hydrogen halide in the presence of an oxidising agent; andeffecting a separation of titanium dioxide obtained from said solution.b) subjecting the aqueous leach solution so obtained to extraction withan immiscible organic phase that selectively leaches iron values intosaid organic solvent.

Yet another aspect of invention provides a method for the separation ofiron values from titanium-bearing ore, comprising the steps of:

a) leaching said ore, or a concentrate thereof, with an aqueous solutionof a hydrogen halide;

b) separating solids from the leach solution obtained in (a), to providea leachate solution;

c) subjecting the leachate solution to extraction with an immiscibleorganic phase that selectively extracts iron values into said organicphase, titanium values in the leachate solution selectively remaining inthe aqueous leachate solution.

d) subjecting the aqueous raffinate so obtained to steps to separateTiO₂ therefrom.

DETAILED DESCRIPTION OF THE INVENTION

Processes for the recovery of titanium dioxide from ilmenite, with highpurity in high yield, are known. Techniques for treating the ilmeniteore, optionally to form concentrate and/or for beneficiation of the oreare known. In some instances, it is possible to treat the ore orconcentrate with concentrated hydrochloric acid solution to effect aleaching of titanium values from the ore or concentrate. In otherinstances, it is necessary or desirable to subject the ore orconcentrate to a smelting step in the presence of carbon and/or fluxingagents, and to then separate a slag from the smelting process which isthen subjected to the leaching step.

One aspect of the present invention is directed to the step of recoveryof titanium values from the leaching solution. The leaching solution isa mixture of aqueous hydrochloric acid containing titanium values, andother soluble material and solid materials, particularly residues of theconcentrate and/or slag from which the titanium values have beenleached. A liquid/solid separation step is conducted, to separate aleachate solution from solids.

The leachate so obtained is treated, according to an aspect of thepresent invention, with an organic phase. The titanium values in theleachate are in the form of titanium halide, especially titaniumtetrahalide which, if hydrochloric acid is used in the leaching step,will be titanium tetrachloride.

In one aspect of the invention, the organic phase is selected so thatiron values are selectively separated into the organic phase. Thus, anorganic/aqueous separation is effected, with the iron values being inthe organic phase and titanium values remaining in the aqueous phase.Preferably, iron values are separated almost to the exclusion of othervalues in the leachate solution, or with values readily separatedtherefrom, so that iron products especially iron oxides, may be obtainedin high purity. Vanadium and other metal values that may be present inthe leachate are preferably retained in the aqueous phase.

Examples of the organic phase are phosphoric, phosphoric acid andphosphinic acids, esters oxides thereof. Specific examples aretri-n-butyl phosphate and di-2-ethylhexyl phosphoric acid. The organicphase preferably contains a diluent e.g. a hydrocarbon, an example ofwhich is a kerosene.

The organic phase may be stripped from the iron values and recycled. Theiron values may then be subjected to pyrohydrolysis, or other steps, torecover iron e.g. as iron oxides. Preferably HCl is obtained as aby-product which is recycled to the step of leaching of ore orconcentrate described above.

In a second aspect, the organic phase is selected such that the titaniumhalide is soluble in the organic phase. Moreover, the organic phase isselected such that the organic phase and titanium halide may beseparated by fractional distillation. The organic phase may be selectedto have a higher or lower boiling point that the titanium halide. Thepreferred titanium halide is titanium tetrachloride. In embodiments ofthe invention, the boiling point of the organic phase differs by least50° C., and especially at least 75° C., from the boiling point of thetitanium halide. The organic phase must be immiscible with the aqueoussolution, such that it forms a second layer so that separation may beeffected.

The titanium halide is extracted from the aqueous solution into theorganic phase, to effect removal of the titanium halide from the aqueoussolution. Such extraction may be carried out in a continuous operationor in a batch operation.

The organic phase may be, for example, a crown ether, phosphine acid,ester or oxide, or tertiary or quarternary ammonium salt.

The organic phase containing the titanium halide is separated from theaqueous solution and from any solid matter, and is then subjected to astep to separate the titanium tetrachloride. In the separation step, theorganic phase containing the titanium halide is heated to effect theseparation of the titanium halide. This is preferably accomplished byvolatilization of the titanium halide, especially if the halide ischloride although for any particular tetrahalide the organic phase maybe selected to effect volatilization of either the titanium tetrahalideor the organic phase. In addition, the organic phase should be selectedso that it has a flash point that is acceptable under the operatingconditions, preferably a flash point above the temperature used inseparation. The organic phase needs to be stable with respect to theaqueous solution and to the titanium tetrahalide at the operatingconditions.

The aqueous solution remaining after extraction may be subjected toknown procedures for recovery of iron or other metal values orprocedures described above, and for recovery and recycle of acid used inthe leaching step. For example, iron oxide (Fe₂O₃) may be recovered andthe acid e.g. HCl, recycled. The organic phase is preferably recycledback to the extraction step, and reused.

The titanium tetrahalide may be subjected to purification steps, ifnecessary. However, if the titanium tetrahalide is volatilized, it maybe of acceptable purity for many end-uses. The titanium halide may beused as such or subjected to further processing steps e.g. to formtitanium metal. Techniques for the conversion of titanium tetrahalideand titanium dioxide to titanium metal are known.

In embodiments of the invention, essentially all of the iron of the feedmaterial i.e. titanium-bearing ore or concentrate is dissolved by theHCl. Thus, sufficient HCl has to be provided. In order to minimise theamount of HCl required in the process, the iron chloride (e.g. H⁺FeCl⁻ ₄or a chloride of iron e.g. FeCl₃) produced in the process issubsequently subjected to pyrohydrolysis to regenerate the HCl. The ironis converted into an iron oxide product (Fe₂O₃). By controlling thecomposition of the iron chloride solution before subjecting it topyrohydrolysis, it is possible to produce a high grade iron oxidesuitable for use in pigment production. If this is not economic, thedisposal of iron oxide becomes an environmental problem. In suchinstances, it is advantageous to remove the iron and upgrade thetitanium ore before subjecting it to treatment with HCl. Thisalternative also has the advantage of decreasing the size of thehydrometallurgical plant for a given production of titanium.

In another aspect of the present invention, the ore or concentrate isleached in aqueous solution in the presence of an acid and an oxidizingagent. A variety of oxidizing agents may be used, including air,hydrogen or other peroxides, or sodium or other perchlorates. Theoxidizing agent should be selected to minimize any contamination of thesolution with cations that have an adverse effect on other processsteps.

In the aqueous solution, the titanium is converted to titanium dioxideand the iron is solubilized. The acid is preferably a hydrogen halide,especially HCl. If the acid is HCl, the concentration of acid may becontrolled, in the presence of the oxidizing agent, to convert iron intoH⁺FeCl⁻ ₄, which is soluble in the aqueous solution. Subsequently,liquid/solid separations may be effected, to separate TiO₂ includingseparation of TiO₂ from tails from the aqueous solution. The aqueoussolution may be treated for recovery of HCl and iron e.g. as Fe₂O₃.

One of the concentration alternatives for the treatment of the ilmeniteore or physically beneficiated ilmenite is the production of atitanium-rich slag and pig iron from it, with the titanium-rich slagbeing subjected to HCl leaching for titanium recovery and the iron beingsold as a foundry-grade pig iron.

The processing of ilmenite to produce titania slag and pig iron isknown. In such processing, the concentrate is calcined in a rotary kilnunder oxidising conditions at about 1200-1300° C. to eliminate sulphurin the ilmenite concentrate. The product is cooled and fed to anelectric furnace with coal/coke as reductant to reduce the iron andproduce a molten slag and pig iron. The electric furnace smelting takesplace at about 1650° C. Disadvantages of the current processing methodinclude (i) the energy in the calcine from the rotary kiln is lost andthe product must be re-heated to the smelting temperature in theelectric furnace with electric power; (ii) the reduction in the electricfurnace produces a lot of CO gas, which often results in the foaming ofthe slag and process control is difficult; and (iii) the reduction ofthe iron oxide by carbon to produce CO is endothermic and the heatrequired is supplied by electricity.

The production of electricity from fuel is typically energy inefficient,to the extent of about 30% conversion, and therefore the process is veryenergy intensive.

In aspects of the present invention, the use of energy is reduced. In apreferred embodiment, the reduction of the calcine is carried out priorto the electric furnace smelting. This may be carried out in a secondreducing section of the calcining kiln or a separate kiln following thecalcining kiln. The reduction is done by reducing gases produced by thepartial reduction of CO (produced in the electric furnace) and/or bypartial combustion of fuel. The reduced calcine so produced istransferred hot from the reduction kiln to the electric furnace, withany additional reductant if required. This saves the energy lost incooling the calcine and improves the efficiency of use of fuel for theiron oxide reduction. The smelting of reduced calcine requires lesselectric energy, produces very little gas in the electric furnace andmakes the electric furnace easier to control. The reduction in thereduction step and the electric furnace are controlled to provide adesired iron level in the slag. This controls the slag meltingtemperature. Depending on the composition of the ilmenite concentrate,the smelting may be carried out at lower temperatures e.g. 1550° C.,thereby conserving more energy compared to the current processing.

Alternatively, the process of smelting and production of moltentitania-rich slag and pig iron in the final step of the above processmay be replaced by a leaching process for removing the reduced ironproduced in the reduction step. This may be carried out by dissolving itin a weakly acidic chloride solution, with aeration, to dissolve theiron and leave a titania-rich oxide product suitable for HCl leaching,as described herein. The advantage of this approach is that thereduction can be carried out at lower temperature e.g. about 800° C.instead of about 1000° C. and the energy of smelting in the electricfurnace is conserved. In addition, in this embodiment, it is possible toeliminate over 95% of the iron and minimise the iron fed to the HClleaching process. In comparison, there is about 90% iron removal in thesmelting route, as some iron has to be left behind to provide a fluidslag in the smelting step. This minimises the usage of HCl and thesubsequent regeneration of the HCl for re-use in the process.

In the production of titanium and TiO₂ pigment, it is necessary toproduce titanium chloride by chlorination of titanium dioxide-containingmaterials e.g. titania slag and rutile concentrate, at about 900° C. inthe presence of coke. By using the present process, the titaniumchloride may be produced by leaching the titania containing materialwith HCl followed by extraction and separation of titanium chloride. Thehigh temperature chlorination is replaced by lower temperatureoperations and avoids the formation of environmentally unacceptabledioxins which can form in the high temperature chlorination. Inaddition, the purity of the titanium chloride produced in the presentprocess will be improved and will require less purification or even nopurification.

In various aspects of the invention, there is provided a process inwhich titanium-bearing ore or concentrate, generally after having beensubjected to a smelting step, is subjected to a leaching step usingaqueous hydrochloric acid. A leach solution containing titanium and ironvalues, and other metallic values depending on the particular ore, isobtained. A liquid/solid separation step is conducted. The solids may besubjected to other separation steps but generally will be gangue.

In preferred aspects of the invention, the leachate solution issubjected to extraction with organic phase to extract iron values, asdescribed above. Such phase may contain 100-200 g/l of iron, or more.Examples of the organic phase are tri-n-butyl phosphate anddi-2-ethylhexyl phosphoric acid, with other examples being describedabove. The organic phase with iron values is then subjected to steps toseparate and recover the organic phase, which is recycled to the step ofextraction of the leachate solution. Iron values, which are in the formof chlorides are recovered, especially by pyrohydrolysis to yield ironoxides and HCl. The HCl is recycled to the leaching of the ore orconcentrate. Iron values of high purity may be obtained.

The leached aqueous solution or raffinate may be treated to separatevanadium and other metallic values, depending on the particular ore,especially by precipitation, to provide a raffinate rich in titanium,and preferably with high-purity titanium values. The titanium values arein the form of the chlorides viz TiCl₄, which may be subjected to stepsto form TiO₂, which is recovered. High purity TiO₂ may be obtained,which may be of sufficiently high purity for use as such.

As an alternative, the raffinate rich in titanium values may besubjected to further extraction, using an organic phase that isimmiscible in water and which has a boiling point that differs from theboiling point of the titanium value, e.g. titanium tetrachloride by anamount to permit fractional distillation. The titanium value isextracted from the aqueous solution of the raffinate into the organicphase and then recovered by distilling or flashing off either theorganic phase or the titanium value, depending on the respective boilingpoints. Titanium metal may then be recovered from the titanium halide.

The by-products of such a process are minimized, and may be treated byknown but relatively simple techniques.

An alternative separation of titanium values is to separate the titaniumas TiO₂ directly from the leaching of the ore or concentrate, byleaching in the presence of an oxidizing agent and separating the TiO₂formed from the solution and from other solids therein.

The present invention provides methods for the separation of titaniumfrom titanium-bearing ores, especially ilmenite. In particular, theinvention provides methods for production of titanium tetrahalides,especially titanium tetrachloride, and TiO₂ with improved purity and/orsuch that related steps in an overall process, including recovery andrecycle of materials, may be simplified and be more cost effective. Inparticular, the volumes of liquid and solids that must be handled in theoverall separation process may be reduced, and associated hardware maybe reduced in size. Such improvements may be of significant economicbenefit.

The present invention is illustrated by the following examples.

EXAMPLE I

A sample of a concentrate of a titanium-bearing ore of a particle sizesuch that 59% by weight would pass through a 100 mesh screen, wassubjected to HCl in an amount estimated to be 100% of the stoichiometricamount of chloride required for the amount of titanium in theconcentrate. The temperature was 95° C.

The amount of concentrated HCl was 541.5 g, for 100 g of concentrate.After a period of 2 hours, 10 g of NaClO₃ were added.

The total treatment time was 3 hours.

The calculated amount of titanium in the samples was 18.5 g, and theamount of iron was calculated to be 39.8 g.

The resultant solution was filtered, washed with an acid solution (HCl)and then with water. The solution were analyzed. The results obtainedwere as follows:

TABLE 1 Vol Wt Assay - mg/l Weight Recovered - g % Recovered Test Sample(ml) (g) Ti Fe Ti Fe Ti Fe Feed Solids 100 19.5 38.7 19.500 38.700Filtrate-Pregnant 400 360 89300 0.144 35.720 0.74 92.30 Acid Wash Water825 17 3560 0.014 2.937 0.07 7.59 Wash Water 500 2.2 100 0.001 0.0500.01 0.13 Tails + 200 Mesh 6.50 9.62 9.68 0.63 0.63 3.21 1.63 Tails −200 Mesh 37.40 47.5 1.15 17.77 0.43 91.10 1.11 % Dissolution 56.10Weight dissolved 56.1 Total 18.549 39.766 95.13 102.76

This example shows the use of oxidant in the recovery of titanium metalvalues, as TiO₂, from a titanium bearing ore. Titanium is effectivelyseparated from iron, as indicated by the high iron content of thefiltrate and the low iron content of the fine tails.

EXAMPLE II

Initial test work on extraction of synthetic titanium chloride solutionshowed that trialkyl phosphine oxide (Cyanex™ 923) could be a goodcandidate for titanium extraction. Tests were then conducted on leachsolution produced by leaching of ilmenite concentrate. These testsshowed that iron preferentially loads into trialkyl phosphine oxide froma solution containing titanium and iron. In a typical test, pregnantleach liquor diluted with an equal volume of distilled water, assaying30.1 g/l Titanium and 30.0 g/l Iron, was contacted with HClpre-conditioned 100% trialkyl phosphine oxide. In this test, sodiumchlorate (NaClO₃), was added to the aqueous phase to raise the EMF ofthe pregnant leach liquor from 406.3 mv to 567.3 mv to aid theextraction. Employing an organic to aqueous ratio of 1:2 and a 30 minutecontact time at 25° C. for extraction yielded an aqueous raffinateassaying 27.5 g/l Titanium and 2.15 g/l iron. This represented anextraction efficiency of 93% for iron and 8.6% for titanium.

This example illustrated that iron could be extracted first from theleach solution.

EXAMPLE III

Test work on different extractants showed that tri-n-butyl phosphate(TBP) is a good candidate for iron extraction from the leach solutionprior to titanium extraction. Large volumes of titanium-containingaqueous leach solution prepared for extraction testing were repeatedlycontacted with 100 vol. % TBP at 40° C. at a 2:1 or 3:1 organic/aqueousphase ratio. After 6 to 8 full contacts, the resulting aqueous raffinateassayed <50 ppm iron. This confirmed the feasibility of extracting ironfrom leach solutions. The titanium is not extracted by TBP and thereforeit is possible to selectively remove iron from leach solution. In thesetests it was seen that when using 100% TBP the iron containing organicis viscous and it was therefore considered preferable to use TBP with adiluent.

A 20% TBP solution in CF-231 kerosene diluent was used to determine theextraction and stripping isotherms. The organic phase waspre-conditioned with 200 g/l hydrochloric acid prior to extraction ofthe iron. The feed aqueous used for these tests analyzed as follows(ppm): Ti 33700, Fe 76000, Al 210, B <5, Ba <1, Ca 150, Cd <5, Co <5, Cr180, Cu 36, K <100, Mg 1450, Mn 160, Mo 5, Na 2400, Ni 4.8, Pb 30, andZn 14.

Varying phase ratios of organic and aqueous phases were tested withbeaker/magnetic stiffer bar contacts at 40° C. for an interval of 10minutes. The observations in these tests were as follows:

1. 20 volume % TBP in CF-231 diluent organic phase saturates at about 22g/l of Fe.

2. The organic-aqueous separation is fast and is complete in about 1minute.

3. A contact time of 10 minutes provides for the transfer from theaqueous to organic phases.

4. These results show that an aqueous pregnant leach solution assaying75 g/l iron, may be extracted using 20 volume percent Tri-ButylPhosphate in CF-231 diluent in two or three stages at 40° C. contacting5 parts Aqueous and l part Organic for 10 minutes.

EXAMPLE IV

After iron extraction, the TBP is stripped to remove the iron and theorganic phase should be returned to the extraction step. Iron extractionwas theoretically shown to require five or six stages using 20 volume %TBP. A stripping isotherm was charted after contacting varying phaseratios of organic and aqueous at 40° C. for 5 minute contact intervals.Leach pregnant solution was used to fully load 20 volume percent TBP in80 volume percent CF-231. Repeated contacts of fresh aqueous with oneorganic volume produced a loaded organic solution assaying about 22 g/liron. This organic solution was stripped with a 13.7 g/l hydrochloricacid solution.

Varying phase ratios of organic and aqueous solution were tested withbeaker/magnetic stiffer bar contacts at 40° C. for an interval of 5minutes. The observations and analytical results are tabled below:

1. Iron loaded organic assaying 22 g/l iron, can be stripped efficientlywith mild HCl at 13.7 g/l concentration in six stages at 40° C.contacting 1 volume of aqueous with 4 volumes of loaded organic.

2. The organic tested viz. 20 volume percent Tri-n-Butyl Phosphate inCF-231 diluent, can be stripped in a six stage counter-current contactto a concentration of 106 ppm iron. This stripped organic can berecycled for extracting iron.

3. By manipulating the concentration of TBP in CF-231 diluent and byincreasing the contact phase ratio to 10:1 organic/aqueous, pregnantstrip solutions can be achieved higher than the 75 g/l Fe obtained inthis isothermal testwork. An example based upon theoretical loading of60 volume % TBP would be an iron concentration of 250 g/L achieved inthe pregnant strip solution.

An example of one run is shown in the attached Table 2.

A bulk TBP raffinate was produced in a series of extraction contacts ofpregnant leach liquor with barren TBP organic. This raffinate analyzedas follows (the composition of the feed aqueous is given in ExampleIII):

Al 500 ppm, B <5 ppm, Ba <1 ppm, Ca 230 ppm, Cd <5 ppm, Co <5 ppm, Cr340 ppm, Cu 44 ppm, Fe <50 ppm, K <100 ppm, Mg 2520 ppm, Mn 390 ppm, Mo<5 ppm, Na 8030 ppm, Ni 16 ppm, Pb 39 ppm, Ti 55600 ppm and Zn 15 ppm.

The stripped iron solution can be used to regenerate the HCl and also itcould be used to produce a value added iron oxide product.

TABLE 2 Solvent Extraction Results Scoping Tests Feed Solution gplDilution A/O Raffinate gpL Extraction Test Solvent Ti Fe Ratio pH RatioTi Fe Ti Fe Comments 13 TBP 3.57 11.7 10:1 1.0 1:1 3:84 5.79 0 50.5 pHadjusted with Ca(OH)₂

EXAMPLE V

The use of diethyl phosphoric acid (DEHPA) with tributyl phosphate (TBP)to remove Vanadium was tested. A larger volume (400 ml) of feed TBPraffinate was treated under sulphur dioxide gas for one hour at 60° C.The pre-reduced aqueous solution was then contacted with HClpre-conditioned TBP/DEHPA organic extractant for 1.5 hours at 60° C.while maintaining the SO₂ bubbling. The organic/aqueous emulsion wasthen allowed to settle and separated. The resulting aqueous raffinatehad evaporated from the feed volume of 400 ml to 286 ml.

A multi-element analysis was carried out to determine vanadium transferinto the solvent. Feed aqueous and resulting TBP/DEHPA raffinate assayswere as follows:

Aqueous Feed TBP/DEHPA Raffinate Al - 560 ppm Al - 1220 ppm Cr - 370 ppmCr - 610 ppm Cu - 48 ppm Cu - 1.5 ppm Fe - 82 ppm Fe - 11 ppm Mg - 2810ppm Mg - 4140 ppm Mn - 460 ppm Mn - 730 ppm Ni - 50 ppm Ni - 34 ppm Ti -31000 ppm Ti - 36500 ppm V - 360 ppm V - 104 ppm Zn - 16 ppm Zn - 1.1ppm

Vanadium was extracted into the aqueous phase at an efficiency of 79.3%in a single extraction. Additionally, iron, nickel, copper and zinc wereloaded into the TBP/DEHPA solvent.

EXAMPLE VI

Precipitation of TiO₂ was also investigated under gas-reducingconditions in conjunction with distilled water to induce titaniumhydrolysis. Tri-n-butyl phosphate raffinate assaying 74.9 g/l Ti and <5ppm Fe was agitated/heated to at least 60° C. while bubbling in 0.5l/min SO₂ gas for a 2 hour interval. Once the temperature was increasedto 95° C., while maintaining SO₂flow, three times the volume of TBPraffinate were added as distilled water. The one-hour reaction period at95° C. yielded a white wet-cake precipitate. Thorough re-pulp washingand oven drying resulted in a light brown solid which assayed asfollows:

Fe   1.28 wt % Cr 0.364 wt. % Si 0.246 wt. % Cu 0.094 wt. % Mg 0.087 wt.% Ni 0.086 wt. % Na 0.052 wt. % V 0.031 wt. % P 0.030 wt. % Zn 0.029 wt.% Mn 0.028 wt. % Pb 0.023 wt. %

All other metals were reported as <0.01 wt. % concentration. Theprecipitation barren solution analyzed only 2.9 ppm titanium, indicatinga precipitation efficiency of greater than 99%. Other metals detected inthe precipitation barren aqueous were follows:

Al 300 ppm Ca 220 ppm Co 4.0 ppm Cr 31 ppm Cu 11 ppm Mg 940 ppm Mn 140ppm Na 70.9 g/L Ni 5.2 ppm V <1 ppm Zn <2 ppm

Example VII

The production of a Titanium Dioxide product of higher purity, at about99 or 99.9% TiO₂, can be achieved by removing all metal contaminantsfrom the leach solution or by selective precipitation. Iron removal hasalready been demonstrated in Example V with the use of tri-n-butylphosphate organic phase.

Titanium 64.7 g/l Iron <50 ppm Aluminum 650 ppm Chromium 260 ppm Copper46 ppm Magnesium 1790 ppm Manganese 310 ppm Nickel 10 ppm Vanadium 47ppm

The more difficult metal contaminants to extract from aqueoushydrochloric media are vanadium and chromium. The use of a reducingagent with the TBP raffinate converts vanadium into thesolvent-extractable V⁴⁺ oxidation state. Chromium may then be separatedfrom titanium by precipitating titanium dioxide at a low pH beforechromium, aluminum, magnesium and manganese are precipitated.

In this example, a small aliquot of TBP raffinate was heated to 70° C.under magnetic stirrer bar agitation. Sulphur dioxide gas was bubbledinto the aqueous solution for 30 minutes. To extract V⁴⁺, an equalvolume of 50/50 TBP(100%)/DEHPA(100%) was added to the aqueous solution.Agitation with SO₂ gas was continued for 1 hour at a temperature of 75°C. This low pH aqueous/organic emulsion qualitatively showed an organicappearance change from a colourless organic to a light golden yelloworganic. Additionally, the heating for an extended period of time causeda pure white titanium dioxide precipitate to separate out of the aqueoussolution.

Although precipitation efficiency of titanium was not established inthis example, the resulting multi-element assay revealed a very pureTiO₂ product. The analytical results were as follows:

Al <0.01 wt. % B <0.01 wt. % Ba <0.01 wt. % Be <0.01 wt. % Ca <0.02 wt.% Cd <0.01 wt. % Co <0.01 wt. % Cr <0.01 wt. % Cu <0.01 wt. % Fe <0.01wt. % K   <1 wt. % Mg <0.01 wt. % Mn <0.01 wt. % Mo <0.01 wt. % Na <0.02wt. % Ni <0.01 wt. % Pb <0.01 wt. % Si <0.02 wt. % V <0.01 wt. % Zn<0.01 wt. %

The resulting precipitate of TiO₂ was oven roasted at 1000° C. to driveoff the hydrated water. The weight loss was 10%. This product obtainedwas >99.0% TiO₂. The composition of the TiO₂ is given in Table 3.

TABLE 3 Composition of TiO₂ TBP/DEHPA Extraction SO₂ Reduction/ppt'nElement Weight % Aluminum Al₂O₃ <0.01 Boron B <0.01 Barium Ba <0.01Beryllium Be <0.01 Calcium CaO <0.01 Cadmium Cd <0.01 Cobalt Co <0.01Chromium Cr <0.01 Copper Cu <0.01 Iron Fe₂O₃ <0.01 Magnesium MgO <0.01Manganese Mn <0.01 Molybdenum Mo <0.01 Phosphorus P₂O₅ <0.01 PotassiumK₂O <0.01 Sodium Na₂O <0.05 Nickel Ni <0.01 Lead Pb <0.01 Silicon SiO₂<0.05 Titanium TiO₂ >99.8 Vanadium V <0.01 Zinc Zn <0.01

To make a direct comparison to the test above, a sample of TBP raffinatewas pH adjusted with 50 weight percent sodium hydroxide to the point ofprecipitating titanium dioxide. The addition of caustic was over a fiveand one-half hour period at an elevated temperature of 40° C. A solidsample obtained was thoroughly washed and dried.

The precipitation barren solution measured at a pH equal to <0.0. Amulti-element analysis of the hydrolyzed product showed metalcontaminants higher than achieved above with DEHPA present. Analysis ofkey elements are as follows:

Chromium <0.01 wt. % Iron <0.01 wt. % Magnesium <0.01 wt. % Manganese<0.01 wt. % Aluminum <0.01 wt. % Vanadium 0.021 wt. % Zinc 0.011 wt. %

Calcination of the precipitate will increase the concentration of TiO₂,but, will also elevate the levels of zinc and vanadium. The use ofTBP/DEHPA allows extraction of vanadium and zinc. Controlled hydrolysisof titanium dioxide at a low pH may result in the prevention ofco-precipitation of chromium.

In another test, an aliquot of TBP raffinate was tested toquantitatively determine the precipitation efficiency of titanium fromTBP raffinate using SO₂ and TBP/DEHPA. The extraction of Vanadium wascarried out at a phase ratio equal to one, but at a lower contacttemperature of 50° C. The organic/aqueous contact interval was extendedto 2 hours at this lower temperature while maintaining the reducingatmosphere with SO₂ gas. After disengaging both phases, the aqueoussolution was separated from the loaded organic. Again, sulphur dioxidegas was bubbled into the aqueous solution while the temperature wasraised to 90° C. or higher. The 150-160 milliliters of treated TBPraffinate was diluted with 300 milliliters of distilled water toinitiate precipitation of the titanium dioxide. The dehydration periodto form TiO2 was extended for 2.5 hours. The resulting white precipitatewas mild HCl washed followed by a distilled water wash and then ovendried at 100° C. The analytical results of the 80 millilitersprecipitation barren and the multi-element analysis of the whiteprecipitate are follows:

Precipitation Barren (80 ml) Washed TiO₂ Precipitate Al 1570 ppm Al<0.01 wt % Cr 1310 ppm Cr <0.01 wt % Cu 4.8 ppm Cu <0.01 wt % Fe 1990ppm Fe <0.01 wt % Mg 4590 ppm Mg <0.01 wt % Mn 790 ppm Mn <0.01 wt % Ni320 ppm Ni <0.01 wt % Mo N/A Mo <0.01 wt % Si <50 ppm Si <0.05 wt % NaN/A Na <0.05 wt % Ti 5590 ppm Ti 58.0 wt % V 1200 ppm V <0.01 wt % Zn 21ppm Zn <0.01 wt %

The analytical results above clearly show that the dehydration oftitanium dioxide takes place at a low pH (pH<0.0), and other metalcontaminants will not co-precipitate. The TiO₂ product obtained is ahydrated oxide that requires further roasting to produce a >99.8% puresolid. The precipitation efficiency of titanium calculates to be 94.8%based on the assay results above.

EXAMPLE VIII

A series of leaching tests were aimed at exploring the phenomenon thatoccurred when leaching at about 90° C. that caused some titanium insolution to precipitate during HCl extraction of ilmenite. At highchloride concentrations and low pH, iron can form anionic chlorocomplexes and halo metallic acid (H⁺FeCl⁻ ₄) and remain in solutionafter leaching. Titanium forms anionic chloro complexes at high chlorideconcentrations and low pH, H₂TiCl₆ with relatively less efficiency. Itis possible to dehydrate dissolved titanium compounds and obtaintitanium dioxide compounds. This approach was tested in a series ofsubsequent tests, to potentially eliminate additional iron reductionwith SO₂ and liquid/solid separation stages. Another additionaladvantage of this approach is the elimination of sulphate in leaching,improving effluent recycling for pyrohydrolysis, potentially eliminatingsulphur in the TiO₂ product and hence in subsequent purification tests.

These tests used a concentrate that contained 19.7% Ti. There were sixtests performed. The first two tests produced 91.7% and 97.6% of the Tiin the precipitate and 98.6% and 96.7% of the Fe in solution. Excess HClwas added and the Ti in the precipitate was decreased. The leaching timewas decreased to 30 minutes from 2 hours and the amount of Ti in theprecipitate was decreased to 80%. In another test, the leachingtemperature was held at 70° C. for two hours then the temperature wasincreased to 95° C. for one hour. The precipitate contained 53.1% Ti.

A further set of leaching tests were aimed at exploring the phenomenonthat occurred when leaching at about 95° C. that caused some titanium toprecipitate with the use of an oxidant. The oxidants tested were oxygengas and sodium chlorate (NaClO₃). Also for these tests, the precipitatewas screened on 200 mesh and the two fractions were assayed for Ti andFe. Screening was used because as a precipitate was formed it would befine material as compared to unleached material (concentrate). For thesetests, the acid level was maintained at 44% excess and oxidant was addedonly after 2 hours of leaching. These tests demonstrated that the Ti canbe precipitated with 95% of the Ti reporting to the minus 200 meshfraction; the iron level was about 1%. The results were reproduciblewith a repeat test and the amount of material treated was increased to200 grams. The amount of iron extracted and present in the leachsolution was about 97% for most of the tests.

Another set of tests was performed to optimize the amount of acid andoxidant used. Lowering the acid used to stoichiometric amount and loweramounts of oxidant caused the amount of Ti reporting to the finefraction to decrease to about 60% and iron level increase to 1.5 to 2%in this fine fraction. If the oxidant was held constant at 1.5 g ofNaClO₃ per 100 grams of concentrate and the acid was increased to 20% or30% excess, the Ti reporting to the fine fraction was 86% with 0.31 to0.51% iron at the respective acid excess level. On the basis of thesetests, an optimum condition is 3g of NaClO₃ per 100 grams of concentrateand 20% excess acid, under which the Ti reporting to the fine fractionis greater than 90% with an Fe level of 0.5% in the fine fraction.Oxygen was tested using the same conditions of acid, time andtemperature and 86% of the Ti reported to the fine fraction and the Felevel was 0.9% in the fine fraction.

A multi-elemental analysis was conducted on the precipitation products,and two results are presented in Table 4. Further details of these testsare provided in Tables 5 and 6.

TABLE 4 Composition of TiO₂ Products made in the Testwork Leach Test ALeach Test B 200 Mesh 200 Mesh Leach Residue Leach Residue ElementWeight % Weight % Aluminum Al₂O₃ 0.32 N/A Boron B <0.01 NA Barium Ba<0.01 N/A Beryllium Be <0.01 N/A Calcium CaO 0.16 1.56 Cadmium Cd <0.01N/A Cobalt Co <0.01 N/A Chromium Cr <0.01 N/A Copper Cu <0.01 N/A IronFe₂O₃ 0.96 0.803 Magnesium MgO 0.33 N/A Manganese Mn <0.01 NA MolybdenumMo <0.01 N/A Phosphorus P₂O₅ N/A N/A Potassium K₂O N/A N/A Sodium Na₂O0.13 N/A Nickel Ni <0.01 N/A Lead Pb <0.01 N/A Silicon SiO₂ 3.74 N/ATitanium TiO₂ >94.35 >97.0 Vanadium V <0.01 N/A Zinc Zn <0.01 N/A

TABLE 5 Test A (110% Stoichiomeric Cl- from HCl used on T12 (+100 M/+150M) Spiral Concentrate) Material 62.0 wt % Passing 100 Mesh Temperature95+ deg C. Time Hours 3 Total Leachant System 413.9 g conc HCl, 5.0 gNaClO₃ oxidant addition Feed Amount 100 grams of T12 (+100/+150 M)Spiral Concentrate Initial Solution 347.8 ml HCl + 0 ml H₂O Assay - VolWt mg/l or wt % Weight Recovered-g % Recovered (Head) Material Balance %Test Sample ml g Ti Fe Ti Fe Ti Fe Ti Fe Feed Solids 100.1 20.42 40.1920.440 40.230 Filtrate- 326 430 110000 0.140 35.860 0.69 89.14 0.7289.94 Pregnant Acid Wash Water 855 68 3360 0.058 2.873 0.28 7.14 0.307.21 Wash Water 510 140 110 0.071 0.056 0.35 0.14 0.37 0.14 1.39 97.27Tails + 200 Mesh 17.30 31.20 5.08 5.40 0.88 26.41 2.18 27.87 2.20 Tails− 200 Mesh 27.40 50.00 0.746 13.70 0.20 67.02 0.51 70.74 0.51 %Dissolution Weight Dissolved 55.4 Totals 19.367 39.872 94.75 99.11100.00 100.00 Calculated Head Titanium 19.3 Iron 39.8

TABLE 6 Test B (120% Stoichiomeric Cl- from HCl used on T12 (+100 M/+150M Spiral Concentrate) Material 63.0 wt % Passing 100 Mesh Temperature95+ deg C. Time Hours 3 Total Leachant System 451.6 g conc HCl, 1.50 gNaClO₃ oxidant addition Feed Amount 100 grams of T12 (+100/+150 M)Spiral Concentrate Initial Solution 379.5 ml HCl + 0 ml H₂O Assay - VolWt mg/l or wt % Weight Recovered-g % Recovered (Head) Material Balance %Test Sample ml g Ti Fe Ti Fe Ti Fe Ti Fe Feed Solids 100 19.5 38.719.500 38.700 Filtrate- 362 1880 101000 0.681 36.562 3.49 94.48 3.4690.83 Pregnant Acid Wash Water 780 64 3480 0.050 2.714 0.26 7.01 0.256.74 Wash Water 505 8.2 110 0.004 0.056 0.02 0.14 0.02 0.14 3.73 97.71Tails + 200 Mesh 10.40 18.40 7.35 1.91 0.76 9.81 1.98 9.73 1.90 Tails −200 Mesh 32.80 51.90 0.480 17.02 0.16 87.30 0.41 86.54 0.39 %Dissolution Weight Dissolved 56.8 Totals 19.67 40.254 100.88 104.01100.00 100.00 Calculated Head Titanium 19.7 Iron 40.3

What is claimed is:
 1. A method for the separation of titanium fromtitanium-bearing ore, comprising the steps of: a) leaching said ore or aconcentrate thereof with an aqueous solution of an hydrogen halide at atemperature of at least 90° C. to provide a leach solution containingtitanium halide; b) separating solids from the leach solution to providea leachate solution; c) subjecting the leachate solution to extractionwith an immiscible organic phase to form an organic phase containingtitanium halide, wherein said organic phase and titanium halide haveboiling points that differ by at least 50° C.; and d) stripping titaniumhalide from said organic phase containing titanium halide by heating tovolatilize titanium halide and effect separation of said titanium halidefrom the organic phase.
 2. The method of claim 1 in which the titaniumbearing ore is concentrated in titanium values by calcining, reductionand smelting in any combination to produce a titanium rich slag andseparate iron as a marketable iron product.
 3. The method of claim 1 inwhich the titanium bearing ore is concentrated in titanium values bycalcining, reduction and leaching to remove iron values.
 4. The methodof claim 1 in which the organic phase is selected from the groupconsisting of crown ether, phosphine acid, phosphine ester, phosphineoxide, tertiary ammonium salt and quaternary ammonium salt.
 5. Themethod of claim 1 in which titanium metal is recovered from the titaniumhalide of step(d).
 6. A method of separating a titanium halide from aconcentrated aqueous solution of the titanium halide, said titaniumhalide being in a concentration such that the titanium halide is stablein said aqueous solution, comprising: a) admixing said aqueous solutionwith an organic phase having a boiling point that differs from theboiling point of the titanium halide by at least 50° C. to form anorganic phase containing titanium halide; b) separating said organicphase containing titanium halide from the aqueous solution; and c)heating the organic phase to strip titanium halide therefrom.
 7. Themethod of claim 6 in which the titanium halide is titanium tetrahalide.8. The method of claim 6 in which the organic phase is selected from thegroup consisting of crown either, phosphine acid, phosphine ester,phosphine oxide, tertiary ammonium salt and quaternary ammonium salt. 9.The method of claim 7 in which the titanium tetrahalide is titaniumtetrachloride.
 10. The method of claim 7 in which the titaniumtetrahalide is separated by volatilization of the titanium tetrahalide.11. The method of claim 7 in which the organic phase has a boiling pointat least 50° C. higher than that of the titanium tetrahalide.
 12. Themethod of claim 7 in which the organic phase has a boiling point atleast 50° C. lower than that of the titanium tetrahalide.